Recovery of copper and zinc as sulfides from copper-iron sulfides

ABSTRACT

A process for recovery of zinc values contained in thermally activated iron containing copper and zinc sulfides in which said activated sulfides have been subjected to an acid leach in aqueous sulfuric acid to produce a solution of ferrous sulfate and a solids residue containing copper and zinc values. The solids residue containing copper and zinc values is reacted with an acidic copper sulfate solution to precipitate copper values as a simple copper sulfide with concurrent dissolution of iron and zinc values to form a solution of ferrous sulfate containing said dissolved zinc values and said solution of ferrous sulfate is separated from the simple copper sulfides and reacted with H 2  S for precipitation of zinc values therefrom as zinc sulfide. Any nickel and cobalt values present in said activated sulfides is dissolved in said solution of ferrous sulfate and reacted with the H 2  S at elevated temperature for selective precipitaton.

This is a division of application Ser. No. 408,132, filed Oct. 19, 1973,now U.S. Pat. 3,964,091.

BACKGROUND OF THE INVENTION

This invention relates to a hydrometallurgical process for treating ironcontaining metal sulfides such as pyritic metal sulfides containingnon-ferrous metal values for separate recovery of said non-ferrous metalvalues and sulfur values wherein said sulfur values can be selectivelyproduced as elemental sulfur or in the form of sulfuric acid to supplyacid requirements of the process.

Many processes are known for the hydrometallurical extraction of desiredmetal values and sulfur from sulfide materials. These hydrometallurgicalprocesses usually comprise the leaching of mineral concentrates atelevated temperatures and pressures in the presence of an oxidant in anacidic medium whereby the non-ferrous metal values can be extracted bydissolution and the sulfur can be removed in elemental form. Seriousdisadvantages of these known processes are the difficulties ofdissolving pyritic copper and iron sulfides and of separatingprecipitated iron compounds from precious metals and dissolvednon-ferrous metals. In most of the known processes, minerals such aschalcopyrite and pyrite are not directly soluble or require such severeleaching conditions that these processes become uneconomical.

Accordingly, it is a principal object of the present invention toprovide a hydrometallurgical process for the treatment of pyritic metalsulfides containing pyrite, chalcopyrite, bornite pyrrhotite and thelike whereby the iron and non-ferrous metal values are rendered activefor selective extraction and separate recovery.

It is another important object of the present invention to provide asubstantially closed and self-supporting hydrometallurgical process fortreatment of the aforesaid pyritic sulfides wherein sulfur values can becontrollably produced as elemental sulfur or in the form of sulfuricacid to satisfy process acid requirements and to substantially obviatethe discharge of sulfur-containing gases to the atmosphere.

A further important object of the invention is the provision of aprocess which permits selective bulk removal of iron values from pyriticmetal sulfides to enable subsequent recovery of non-ferrous metalsulfides in enriched and activated form with attendant economies inoperating and capital costs.

Another important object of the present invention is the provision of aprocess for the treatment of pyritic copper sulfides for the recovery ofcopper values and precious metals together with metal sulfides such aszinc, cobalt and nickel sulfide and molybdenite.

SUMMARY OF THE INVENTION

Accordingly, the process of the invention may be broadly described asfollows. Ores and concentrates of pyritic metal sulfides are thermallyactivated by sequentially heating and reducing said sulfides in acountercurrent flow of a substantially non-oxidizing gas and a reducinggas respectively in a reactor whereby the reaction products of thereducing gas and sulfides together with liberated labile sulfur arecontrollably combusted with oxygen to satisfy heat requirements of thethermal activation process and to convert sulfur values to SO₂ gas. Themajor portion of the sulfidic iron present in the sulfides is convertedto an acid leachable form and non-ferrous metal values are activated andrendered amenable to an acid oxidation leach. The thus activatedsulfides are subjected to an acid leach in aqueous sulfuric acid forproduction of ferrous sulfate and evolution of H₂ S gas which areseparated from the solid sulfide residue containing concentratednon-ferrous metal values. The ferrous sulfate is oxidized and hydrolysedto precipitate the iron, which is removed from the system, and toregenerate sulfuric acid which is recycled to the acid leach. At least aportion of the SO₂ gas from said heating step and at least a portion ofthe H₂ S from the acid leach are converted to sulfuric acid to satisfythe acid requirements of the process, and remaining SO₂ is combined andcatalytically reacted with remaining H₂ S to produce elemental sulfur.

The solids residue from the acid leach, containing enriched copper andother non-ferrous metal values, can be subjected directly to an acidoxidation leach in aqueous sulfuric acid in which copper values arecarried into solution as a pregnant copper sulfate liquor which ispurified and is subjected to recovery of copper by electrowinning orhydrogen reduction. Remaining oxidation leach residue can be furtherprocessed to yield elemental sulfur and a concentrate containingprecious metals and any molybdenum present in the feed sulfides.

Alternatively, the solids residue from the acid leach can be subjectedto a wet activation leach with copper sulfate, prior to the acidoxidation leach, for further and selective activation of copper valuesand for dissolution of iron values and certain non-ferrous metal values,such as zinc, nickel and cobalt, for their subsequent recovery.

BRIEF DESCRIPTION OF THE DRAWING

The process will now be described with reference to the accompanyingflowsheet which schematically illustrates the steps thereof.

DESCRIPTION OF THE PREFERRED EMBODIMENT

Ores and concentrates suitable for treatment for recovery of non-ferrousmetal values can be low grade pyritic copper sulfides, such as low gradeconcentrates of bornite, chalcopyrite, and mixtures thereof withpyrrhotite and pyrite, and can contain precious metals and other metalvalues such as zinc, cobalt, nickel and lead sulfides and molybdenite,as well as small amounts of arsenic, antimony, selenium, tellurium andtin.

Particle size of concentrate has been found to be not critical forsuccessful processing, the normal degree of comminution necessary forbeneficiation usually being satisfactory for rapid activation andleaching.

The concentrate is fed to the thermal activation step 1 which consistsof a two-stage heating and reduction operation preferably carried out ina countercurrent reactor such as, for example, a vertical multiplehearth roaster. The concentrate is fed to a heating zone in the top ofthe reactor and travels downwardly through the hearths countercurrent toan ascending flow of a hot inert or substantially non-oxidizing gas, tobe described hereinbelow, wherein the temperature of the concentrate israised to a temperature in the range of from about 550° C. to atemperature not exceeding the fusion point of the concentrate,preferably from about 650° to about 800° C. Heating of the copperconcentrate to this temperature range results in thermal decompositionof the pyritic sulfides with the liberation of labile sulfur. Thereactions may, for example, be expressed as follows:

    5 CuFeS.sub.2 → 5 CuFeS.sub.1.8 + S.sup.0           (1)

    feS.sub.2 → FeS.sub.1.05 + 0.95 S.sup.0             (2)

it is understood that the exact composition, and particularly the sulfurcontent of the delabilized sulfides, may vary and is dependent onprocess conditions.

The delabilized sulfides then descend through a reducing zone of theroaster having a countercurrent ascending flow of a reducing gas thereincapable of accepting sulfur, such as hydrogen, or a hydrogen containinggas such as, for example, ammonia synthesis gas, reformed natural gas ormethane, which reacts with the delabilized sulfides at temperatureswithin the range of from about 550° C. to a temperature not exceedingthe fusion point of the concentrate, preferably from about 650° to 800°C. to form a bornite-pyrrhotite type mixture, as follows:

    5CuFeS.sub.1.8 + H.sub.2 → Cu.sub.5 FeS.sub.4 + 4 FeS + H.sub.2 S (3)

    feS.sub.1.05 + 0.05 H.sub.2 → FeS + 0.05 H.sub.2 S  (4)

the copper iron sulfide formed according to reaction (3) has theapproximate empirical formula of bornite but its actual compositionresembles a mixture of a bornite type material and ferrous and cuproussulfide. The pyrrhotite formed according to reactions (3) and (4) isreadily acid leachable.

We have found that adequate utilization of the hydrogen in the reductionzone can be achieved by countercurrent contacting of the hydrogen withthe delabilized sulfides.

The H₂ S produced rises into the aforesaid heating zone for mixture withair or an oxygen containing gas in an amount only sufficient, i.e. asubstantially stoichiometric amount of oxygen, for controlled combustionof the H₂ S, H₂ and labile sulfur to form SO₂ and H₂ O which aredischarged from the heating zone, and to supply heat requirements forthe heating zone. H₂ S from a subsequent acid leach may be introduced ifdesired as indicated by the broken line 18 in the drawing to the heatingzone to supplement heat requirements when processing low-pyrite metalsulfides. In case of excess heat the temperature of the solids may becontrolled by injection of water, by removal of combustible gases, suchas H₂, H₂ S and sulfur vapor, or by injecting air at points close to thegas discharge end of the reactor.

It is important that the solids in the heating zone be maintained in asubstantially neutral or partially reducing state to prevent theformation of metal oxides.

The atmosphere in the heating and delabilizing zone is maintained onlysufficiently oxidizing by the addition of, for example, only enough airor an oxygen containing gas to permit burning of the combustible gasesand labile sulfur, and to heat the concentrate to the desireddelabilizing and reducing temperatures. The atmosphere in the heatingzone is defined as preferably being a substantially neutral or partiallyreducing atmosphere in relation to the solids being treated andcontaining less than 5% by volume oxygen.

It is important in the heating and reducing zones that the concentratenot be heated to a temperature at which phenomena such as sintering andbridging may occur. By agglomerating the feed prior to introduction tothe thermal activation step, such as for example by pelletizing orbriquetting the feed, thermal activation can be carried out at anelevated temperature while minimizing sintering and bridging. We havefound that a close temperature control can be achieved and maintained byadding controlled amounts of air at spaced intervals along the length ofthe heating zone of the reactor. For example, a concentrate temperatureof about 700° C. can be established by selectively adding air or otheroxygen containing gas along the heating zone such that the gaseousatmosphere is heated to a temperature not exceeding 760° C. Utilizationof hydrogen gas and acid solubility of resultant iron values decrease atlower temperatures and thus it is desired to maintain the concentratetemperature, during the thermal activation, as high as possible withoutoccurrence of the above named phenomena. Some of the As, Sb and Sepresent in the concentrate is volatilized in the thermal activationstep.

A typical pelletized copper concentrate, analyzing 23.4% copper, 30.2%iron, 4.8% zinc and 35.0% sulfur, was continuously processed in amultiple hearth roaster by heating of the concentrate in a upper heatingzone of the roaster by the controlled combustion of hydrogen and gaseousreaction products described above with air to achieve hearthtemperatures ranging from 640° to 735° C. The heated concentratedescended through a hydrogen atmosphere for a total retention time of 30minutes. Samples of the discharge solids were subjected to acidleaching, to be described below, resulting in iron extractions thatincreased with higher temperatures, as shown in Table 1.

                  TABLE 1                                                         ______________________________________                                                  Average Hearth   Fe                                                 Test      Temperature ° C.                                                                        Extraction %                                       ______________________________________                                        1.        640              56                                                 2.        700              63                                                 3.        705              73                                                 4.        735              83                                                 ______________________________________                                    

The thermally activated concentrate is fed to an acid leach 2 where itis mixed with a solution containing from about 60 to 300 gpl freesulfuric acid, preferably in an amount of up to about 20% on excess ofstoichiometric requirements to convert iron to FeSO₄, at a temperatureof from about ambient to about 200° C. Preferably the leach is conductedat a temperature of from about 30° to 100° C. at atmospheric pressure,for a time sufficient to leach from about 60 to about 90% iron as FeSO₄,as represented by the following general equation:

    Cu.sub.5 FeS.sub.4 + FeS + H.sub.2 SO.sub.4 → Cu.sub.5 FeS.sub.4 + FeSO.sub.4 + H.sub.2 S                                    (5)

the free acid content of the final solution drops to about 15 gpl,usually from an initial content of about 130 gpl, before leaching ofiron ceases, to produce a solution containing from about 50 to 90 gpliron as FeSO₄. This bulk removal of iron leaves an iron depleted leachresidue containing copper and iron in a molar ratio of at least about 1: 1, usually in the range of from about 2 : 1 to 15 : 1, which issuitable for an acid oxidation leach. Temperature is not critical withinthe preferred range once the reaction has been initiated, a temperatureof about 30° to 40° C. usually being required to start the reaction,with temperatures in excess of 100° C. decreasing the solubility of theFeSO₄.

To control foaming, the reaction can be moderated by slow acid additionor by the use of two-stage countercurrent leaching. If necessary,chemical or mechanical foam breakers may be used. Two-stagecountercurrent leaching is also preferred in that iron extraction isenhanced and the FeSO₄ liquor has low acidity which is desirable for thesubsequent iron precipitation step. The first stage is conducted infiltrate from the second stage containing from about 60 to about 150 gplfree sulfuric acid and about 60 gpl Fe for a time sufficient to obtainan iron extraction of from about 60 to 70% of the acid leachable ironwith iron content in the filtrate of about 80 to 90 gpl and free acid ofabout 1 to 5 gpl. The solids residue from the first stage is charged tothe second stage where it is contacted with sulfuric acid generated inother steps of the process, to be described, containing in excess of 200gpl free sulfuric acid. The slurry is reacted for a time sufficient toextract the remainder of the acid leachable iron and the filtrate isrecycled to the first stage leach.

The filtrate from the first stage is fed to an iron removal and acidregeneration step. The residue from the second stage is suitable for asubsequent acid oxidation leach, or if necessary, for a wet activationleaching step prior to the acid oxidation leach to remove additionaliron and to provide a copper free solution suitable for the removal ofzinc to control zinc build-up in the system.

Samples of roaster discharge solids thermally activated as describedabove were repulped with water and concentrated sulfuric acid,equivalent to a H₂ SO₄ : Fe molar ratio of 1.8 : 1.0, was addedstagewise to simulate two-stage atmospheric leaches extending over aperiod of 1 hour. Temperature was maintained at about 100° C. Ironextractions up to 81.3% were obtained with copper to iron molar ratiosin the residue of about 4 : 1.

Hydrogen sulfide from the acid leach reaction and SO₂ in the exit gasesfrom the heating zone of the thermal activation are utilized in varioussteps of the process with excess being converted to elemental sulfur.Portions of the H₂ S and SO₂ are fed to an acid plant 3, well known inthe art, for the production of sulfuric acid, after combustion of the H₂S, in an amount sufficient to satisfy the acid requirements of theprocess. Excess H₂ S and SO₂ are fed to a Claus plant 4, also well knownin the art, for catalytic reaction and production of elemental sulfurwhich is stable and relatively inert and thus suitable for storage orshipment. Unconverted H₂ S from the Claus plant is combusted to SO₂ andfed to the acid plant. A small quantity of SO₂ is used in the reductionof ferric iron, to be described and small quantities of H₂ S may be usedin the thermal activation and in the precipitation of zinc, cobalt andnickel, as required.

The acid leach liquor containing from 50 to 90 gpl iron as FeSO₄ andfrom 5 to 60 gpl free H₂ SO₄ is fed to the oxydrolysis step 5 where ironis precipitated from the solution in a one-stage oxidation andhydrolysis operation and sulfuric acid is produced. Preferably ironcompounds of the jarosite type are precipitated. Ammonia or a suitableammonium or alkali-metal compound is added to the leach liquor in anamount of from about 50 to 120%, preferably from about 80 to about 100%,of stoichiometric requirements, together with an oxygen bearing gaswhich provides an oxygen partial pressure in the range of from about 50to about 200 psi, preferably about 150 psi. A temperature above about140° C., preferably in the range of from about 140° to about 230° C., ismaintained and the oxydrolysis of iron in the presence of cations suchas ammonium, sodium or potassium is permitted to proceed for 30 to about90 minutes to produce jarosite ( [A Fe₃ (SO₄)₂ (OH)₆ ] where A = NH₄ ⁺,Na⁺ or K⁺) and a dilute sulfuric acid. The sulfuric acid solutioncontaining some residual iron is recycled to the acid leach step afterreduction of ferric iron to the ferrous state with SO₂ in step 6. Thereduction is carried out according to well known methods at atemperature of about 70° to 80° C. under atmospheric pressure.

A leach liquor, containing 3.7 gpl copper, 51.3 gpl iron and 23.0 gplfree sulfuric acid, to which had been added 90% of the stoichiometricquantity of NH₄ ⁺ ion (as NH₄ OH) required to form ammonium jarosite,was subjected to oxydrolysis by heating with agitation in a pressurevessel at a temperature maintained at 180° C. under 180 psi of oxygenpartial pressure. The mean residence time of solution in the pressurevessel was 42 minutes. The filtrate of the resulting product analyzed2.5 gpl copper, 3.8 gpl iron, and 36.2 gpl sulfuric acid. The solidsanalyzed <0.1% copper, 34.7% iron and 13.1% total sulfur.

The thermal activation and acid leach steps, with integrated productionof elemental sulfur and sulfuric acid, and oxydrolysis of the FeSO₄liquor for production of sulfuric acid removal of the major portion ofthe iron from the system, form an essentially closed system which can bereadily balanced to supply process acid and heat requirements withexcess sulfur produced in a stable elemental form without the dischargeof noxious fumes and gases to the atmosphere.

The solids residue from the acid leach can be directly fed to an acidoxidation leach 7, or optionally fed to a wet activation leach step 8shown by broken lines in the drawing, if the acid leach residuescontain; beside copper values, significant amounts of zinc, lead, nickeland cobalt values and residual iron values. In leach 8, the total amountof iron which can be selectively removed from the process can beincreased significantly, the recovery of precious metals from the leachresidue can be facilitated due to iron removal, the copper values arefurther concentrated, and a control over the build-up of zinc in theprocess, which is the most common attendant metal, can be exercised byprecipitation of dissolved zinc in a subsequent step, as will becomeapparent as the description proceeds.

In the wet activation leach 8, the solids residue from the acid leach isreacted with an acidic copper sulfate solution for replacement by copperof residual iron and any zinc, lead, nickel and cobalt values present inthe said solids residue with production of activated copper sulfides,lead sulfate and a solution of ferrous sulfate and a sulfate of any ofsaid zinc, nickel and cobalt values present. For example, iron in thebornite-like material is replaced by copper from the acid CuSO₄ solutionaccording to the following equation:

    Cu.sub.5 FeS.sub.4 + CuSO.sub.4 → 2 Cu.sub.2 S + 2 CuS + FeSO.sub.4 (6)

iron is removed from the acid leach residue down to a level of from 2 to10% Fe in the activated copper sulfides, at a copper utilizationcorresponding to a Cu (precipitated)/Fe (solubilized) molar ratio ofclose to unity.

The wet activation leach is conducted at a temperature within the rangeof from about 140° to 200° C., preferably about 160° C., underautogenous pressure, and at a pH of about 0.5 to 2, with a portion ofthe acid CuSO₄ solution from the acid oxidation leach step 7. The acidCuSO₄ solution contains from about 20 to 100 gpl Cu, from about 3 to 15gpl Fe, from about 5 to 35 gpl H₂ SO₄, and about 20 gpl Zn.

Ferric iron in the feed solution to the wet activation leach may lead tothe formation of elemental sulfur as exemplified by the followingequation:

    Fe.sub.2 (SO.sub.4).sub.3 + CuS → CuSO.sub.4 + 2 FeSO.sub.4 + S.sup.0                                                   (7)

the elemental sulfur reacts further to produce sulfuric acid asexemplified by one of several reactions:

    S.sup.0 + 3 Fe.sub.2 (SO.sub.4).sub.3 + 4 H.sub.2 O → 6 FeSO.sub.4 + 4 H.sub.2 SO.sub.4                                        (8)

the copper sulfate formed according to (7) must be subsequentlyrecovered and the sulfuric acid generated according to (8) undesirablyadds to the acidity of the system. The amount of ferric iron fed to thewet activation leach therefore should be kept to a minimum to avoid boththe dissolution of copper and the production of excess acid.

The reactions involving Zn-, Pb-, Ni- and Co-sulfides also occur asfollows:

    MeS + CuSO.sub.4 → CuS + MeSO.sub.4                 (9)

where Me = Zn, Pb, Co, or Ni. These reactions, which consume an amountof CuSO₄ dependent on the amount of these sulfides present, provide aneffective means for dissolution and subsequent recovery or removal ofZn, Co and Ni from the system while Pb precipitates as its sulfate. Forexample, copper concentrates often contain sphalerite which is at leastpartly leached in the wet activation leach 8.

Solids residue from the preceding acid leach was mixed with an acidiccopper sulfate solution containing 70 gpl copper as copper sulfate andagitated in a closed vessel for 1 hour at a temperature of 160° C. Theamount of copper provided in solution was equivalent to a copper to ironmolar ratio of 1.04 : 1, or copper to iron + zinc molar ratio of 0.85: 1. The solids product separated from the solution analysed 55.7%copper, 7.15% iron, 2.58% zinc and 23.9% total sulfur; indicating ironand zinc extractions from the solids residue of 54.7% and 39.9%respectively. The combined iron extraction in the acid and the wetactivation leaches was 87.2% and the combined zinc extraction from theseleaches was 70%.

The solution from the wet activation leach 8 is a substantiallycopper-free FeSO₄ liquor which is subsequently stripped of zinc in zincprecipitation 9 wherein a partial pressure of H₂ S of from about 30 toabout 100 psi, preferably about 100 psi, at a temperature of fromambient to about 100° C., is maintained to precipitate the zinc as aconcentrate of ZnS. Zinc removal from solution decreases with increasingacidity. A temperature of about 70° C., a normal operating temperature,and acidities provided by free acid contents in the range of from about20 to 40 gpl sulfuric acid, are acceptable for zinc removal. Theremaining acid FeSO₄ liquor, containing less than about 2 gpl Zn, isreturned to the acid leach to pick up more iron as FeSO₄, oralternatively is fed to oxydrolysis 5 for removal of iron, as indicatedby the broken line 20 in the drawing. Cobalt and nickel in the solutionfrom the wet activation leach 8 do not precipitate with H₂ S under theconditions maintained in zinc precipitation 9, hence, substantiallycomplete separation is possible. If desired, nickel and cobalt can beprecipitated from the residual solution from the zinc precipitation byraising the temperature of the solution to as high as 150° C.,preferably in the range of from about 100° to 150° C., and maintaining apartial pressure of H₂ S of about 100 psi. A concentrate of cobalt andnickel sulfide is recovered for further processing.

The operation of the zinc precipitation was typified by treating arepresentative solution from the wet activation leach, containing 0.019gpl copper, 41.3 gpl iron, 13 gpl zinc and 58 gpl sulfuric acid, byagitation in a pressure vessel for 30 minutes at 70° C. under 100 psipressure of H₂ S gas for precipitation of zinc sulfide. The resultingslurry was filtered to yield a filtrate which contained 1.1 gpl zinc.The solids analyzed 60.8% zinc, 32.4% sulfur, 0.15% copper and 0.15%iron.

The activated copper sulfides from the wet activation leach 8, or thesolids residue from acid leach 2, is fed to a single or multi-stage acidoxidation leach 7 and leached with return acid from copper recovery 16according to the following equation: ##EQU1## in which x, y and z havevalues in combinations covering various copper sulfides and copper-ironsulfides. For example, for Cu₂ S, x has the value 2, y has the value 0and z has the value 1. Under oxidizing conditions the ferrous sulfateusually is at least partially oxidized to ferric sulfate. The mainreaction taking place in the oxidizing leach is the dissolution ofcopper as cupric ion and the oxidation of sulfide sulfur to elementalsulfur. Sulfides such as those of silver and molybdenum are relativelyunreactive and, together with precious metals, substantially remain assolids in the leach residue together with the gangue materials andelemental sulfur. Zinc and lead sulfides will react similarly to thecopper to form ZnSO₄ and PbSO₄. The acid for this leach is return acidfrom a subsequent copper recovery step containing from about 40 to 200gpl H₂ SO₄, from about 0 to 40 gpl cupric ions and from about 0 to 10gpl ferrous ion. The composition of the return acid will depend largelyon the method of copper recovery used. If, for example, hydrogenreduction is employed, the return acid may contain from about 80 to 120gpl H₂ SO₄, from about 10 to 25 gpl copper as cupric ions and from about1 to 5 gpl iron as ferrous ions.

To achieve high rates of reaction, the temperature should be controlledjust below the melting point of elemental sulfur. The upper limitingtemperature thus is that temperature at which the elemental sulfurmelts, i.e. 113° to 119° C., causing occlusion of unreacted sulfides andthereby inhibiting the leaching reaction. We have found in two-stagecountercurrent oxidation leaching that by maintaining the temperature ina first stage leach in the lower end of the range, i.e., within therange of from about 70° to 100° C., the blinding of particle surfaces iseffectively prevented, while the temperature in the second stage of theleach may be maintained in the higher end of the range, i.e., within therange of from about 105° to about 110° C.

The oxidation leach can be performed under a wide range of pressures. Wehave found that in the presence of an oxygen bearing gas the pressure inthe oxidation leach may be up to about 400 psi partial pressure ofoxygen to provide satisfactory extractions. Pressures above 400 psiimprove the leaching rates marginally and do not justify the additionalcapital costs for higher pressure equipment. Thus, it is possible toconduct the oxidation leach with an oxygen bearing gas providing apartial pressure of oxygen in the range of from about atmospheric toabout 400 psi, preferably at about 200 psi. An effective oxidation leachcan be obtained by maintaining the pressure in a first stage of theoxidation leach in the lower end of the above stated pressure range atfrom substantially atmospheric to about 300 psi, and to maintain thepressure in the second stage in the higher end of the range of fromabout 100 to about 400 psi.

The oxygen bearing gas can be air, oxygen, or mixtures thereof,preferably oxygen. We have found that a substantial portion of thecopper sulfides such as Cu₂ S, CuS and Cu₅ FeS₄ will be extracted in thepresence of the oxygen bearing gas in the first stage oxidation leachwhile remaining copper sulfides will be extracted in the second stageleach.

The retention time of the reaction mixture in each of the stages of theoxidation leach may vary within a broad range and is dependent not onlyon the required extraction in each stage but also on the conditions ineach stage. Thus we have found that at higher temperatures and pressuresthe retention time may be between 10 minutes to 30 minutes while atlower temperatures and pressures the retention time may vary between 1hour and 12 hours.

The acidity of the reaction mixture in the oxidation leach is criticaland must be carefully controlled to ensure a high dissolution of thecopper as well as to prevent the precipitation of iron compounds andassociated loss of silver and molybdenum. If the pH of the reactionmixture rises too high, the iron in the solution will hydrolyse andprecipitate, and valuable silver and molybdenum which are present in thereaction mixture will be occluded by this precipitate and will leave theprocess in the tailings from the subsequent flotatiion step. The acidityis controlled by the addition of return acid, or sulfuric acid from theacid plant, to prevent the precipitation of iron.

An oxidation leach was conducted as follows. Solids residue from thepreceding wet activation leach containing 53.4% copper, 11.6% iron,2.14% zinc and 24.8% total sulfur were repulped with an electrolytecontaining excess sulfuric acid such that the sulfuric acid to copper +zinc + iron molar ratio was 1.05. The slurry was agitated in a pressurevessel with an oxygen partial pressure of 200 psi at a temperatureinitially maintained at 95° C. and raised to 110° C. during the latterportion of the acid oxidation leach. Table 2 indicates analyses of thesolid and liquid constituents of samples taken at 1 hour intervals fromthe commencement of feed.

                                      TABLE 2                                     __________________________________________________________________________    Time                                                                              Filtrate (gpl)                                                                             Residues (%)  Extractions (%)                                (Hrs.)                                                                            Cu Zn Fe H.sub.2 SO.sub.4                                                                  Cu  Zn Fe  S°                                                                        Cu Zn Fe                                       __________________________________________________________________________    Feed                                                                              20 20 5  145 53.4                                                                              2.14                                                                              11.6                                                                             24.8                                              1   75.4                                                                             21.4                                                                             10.1                                                                             43.8                                                                              5.61                                                                              1.69                                                                             9.6 37.8                                                                             95.6                                                                             67.0                                                                             65.3                                     2   80.8                                                                             22.4                                                                             15.6                                                                             24.0                                                                              2.23                                                                              0.73                                                                             3.9 61.5                                                                             98.6                                                                             88.3                                                                             88.4                                     3   83.4                                                                             23.4                                                                             18.3                                                                             16.0                                                                              0.99                                                                              0.17                                                                             1.0 70.3                                                                             99.4                                                                             97.5                                                                             97.2                                     4   82.0                                                                             22.8                                                                             18.0                                                                             13.9                                                                              0.79                                                                              0.08                                                                             0.7 67.3                                                                             99.5                                                                             98.8                                                                             97.9                                     __________________________________________________________________________

The solids in the discharge slurry from the oxidation leach, containingelemental sulfur, unreacted sulfides, lead sulfate, molybdenum andprecious metals, are separated from the leach copper liquid and aretreated for recovery of these values. Dependent on the quantity ofgangue materials in the discharge slurry, the slurry may be subjected,if desired, to flotation (not shown) to separate the insoluble ganguematerial from a concentrate which is treated for recovery of the abovenamed values.

The discharge slurry, or the concentrate from flotation is subjected toa pelletization step 10. Pelletizing of the elemental sulfur is effectedby heating the slurry to about 125° C. under constant agitation for atime sufficient to permit the elemental sulfur to melt and coalesce andthen cooling to about 90° C. for a time to enable the pellets tosolidify. These pellets are separated by standard techniques. Forinstance the resulting slurry is screened on a 35 mesh Tyler screen andthe +35 mesh pellets washed to remove slimes. The -35 mesh pellets andslimes are subjected to centrifugal separation in a hydrocyclone. The+35 mesh pellets and hydrocyclone underflow are heated to from 120° to140° C. and subjected to hot filtration 11 at a pressure of about 20psig air pressure to produce an elemental sulfur product, whichcomprises from 80 to 95% of the sulfur present in the discharge slurryfrom the oxidation leach.

The slimes and filtercake from filtration 11, consisting of a smallamount of unreacted copper sulfides, hydrolysed iron, sulfur and gangueminerals, together with precious metals, tin, lead sulfate andmolybdenum, can be processed for further metal recovery by severaltechniques. For example, the slurry can be subjected to an oxidizingacid leach 12 at a temperature in the range of from about 200° to 250°C. and an oxygen partial pressure in the range of from about 150 to 250psi to dissolve copper values as CuSO₄ and iron as FeSO₄, in which thelatter oxidizes and hydrolyses to basic ferric sulfates or hydratedoxides. These iron products can be redissolved, if desired, at 90° C.with an SO₂ treatment (not shown) leaving a final residue containingprecious metals, tin, lead sulfate and molybdenum. Any silver that issolubilized can be recovered by cementation with copper powder in asilver sponging step 13. The remaining liquor containing copper and ironvalues and sulfuric acid, together with residual acid from thesubsequent copper recovery, to be described, is recycled to theoxidation leach 7.

The pregnant copper liquor from the acid oxidation leach 7, containingabout 60 to 100 gpl copper, about 5 to 25 gpl sulfuric acid and about 3to 15 gpl ferric ion, is purified by a two-stage process, preparatory torecovery of copper values, by removal of major portions of silver,selenium, tellurium, arsenic, antimony, tin and iron. Silver is firstprecipitated from the pregnant liquor by the addition of copper powderin silver cementation 14. Copper powder is added in an amount of about10% in excess of stoichiometric requirements to reduce ferric iron toferrous iron and at a temperature in the range of from about ambient toabout 100° C., preferably from 60° to 70° C., whereby the silver contentis reduced to less than 5 mgpl, some selenium is removed, and the ironis reduced to the ferrous state. The liquor is treated in purification15 by heating to from about 150° to 250° C., preferably to about 200°C., under an oxygen partial pressure of from about 20 to about 200 psi,preferably about 150 psi, to precipitate iron as ferric oxide and basicsulfate and to remove most arsenic, antimony, tellurium and selenium. Itis important that the pregnant liquor does not contain more than 10 gplfree acid, corresponding to a pH of about 1-2. Any excess free acidtherefore is neutralized with CaO which can be injected directly intothe purification reactor.

The two-stage purification is exemplified by the following example.Solution from the oxidation leach, containing 69 gpl copper, 4.9 gpliron (1.2 gpl Fe⁺ ⁺ ⁺ and 3.7 gpl Fe ⁺ ⁺), 21.4 gpl sulfuric acid, 1.0mgpl silver and 0.38 mgpl selenium, was mixed with 6 gpl copper dust,stirred in the absence of air for 20 minutes at 60° to 70° C., and thenfiltered through a bed of copper powder. The filtrate analyzed 67 gplcopper, 4.4 gpl iron (0.1 gpl Fe⁺ ⁺ ⁺ and 4.3 gpl Fe⁺ ⁺), 19 gplsulfuric acid, 0.2 mgpl silver and 0.25 mgpl selenium. The substantiallysilver-free solution was further processed by first neutralizing thesulfuric acid to 10 gpl by the addition of lime, filtered, and thenagitated in a pressure vessel at a temperature maintained at 200° C.under a partial pressure of oxygen of 150 psi for a retention time of 35minutes. The resulting slurry was filtered to yield a solids residuewhich contained 2.1% copper, 52.6% iron and 4.6% total sulfur. Thefiltrate analysed 66 gpl copper, 0.5 gpl iron, 17.6 gpl sulfuric acid,0.3 mgpl silver and 0.12 mgpl selenium. The filtrate substantially freeof iron, silver and selenium, was suitable for subsequent copperrecovery.

The iron oxide slurry is thickened, filtered, washed free of copper andfed to oxydrolysis step 5. The pregnant liquor, essentially a pure CuSO₄solution containing less than 1 gpl iron, can be fed to copper recovery16 for recovery of metallic copper by electrowinning or by hydrogenreduction.

Pure copper cathodes can be recovered from the purified CuSO₄ pregnantliquor in a standard electrowinning tank house. The copper containingliquor is diluted with a recycle of spent electrolyte for a cell liquorfeed containing about 45 gpl Cu and 100 gpl H₂ SO₄ and fed toelectrolytic cell for production of cathodes at a current density in therange of from about 180 to 400 amperes per square meter and a celldischarge liquor of 20 gpl Cu and about 140 gpl H₂ SO₄. Spentelectrolyte is recycled to the oxidation leach step 7.

Alternatively, copper can be won from the CuSO₄ pregnant solution byhydrogen reduction. We have found that a direct reduction with hydrogengas at elevated temperature and pressure in a continuous fashion is veryrapid. The reaction is carried out on a continuous basis in a suitablereaction vessel with agitation at a temperature of about 130° to 200° C.under a nitrogen partial pressure of about 300 to 400 psi. The retentiontime required for the reduction of copper from about 80 gpl to thetarget residual amount of 20 gpl is not more than about 30 minutes.Residual solution is recycled to the oxidation leach step 7.

What we claim as new and desire to protect by letters patent of theUnited States is:
 1. A process for recovery of zinc values contained inthermally activated iron containing copper and zinc sulfides in whichsaid activated sulfides have been subjected to an acid leach in aqueoussulfuric acid to produce a solution of ferrous sulfate and a solidsresidue containing copper and zinc and separation of said solutioncontaining ferrous sulfate from said solids residue which comprises:reacting said solids residue containing copper and zinc with an acidiccopper sulfate solution containing from about 20 to about 100 gpl copperand from about 5 to about 35 gpl sulfuric acid at a temperature in therange of from about 140° to about 200° C. under autogenous pressurewhereby copper values are precipitated as a simple copper sulfide withconcurrent dissolution of iron and zinc values to form a solution offerrous sulfate containing said dissolved zinc values, separating saidsolution of ferrous sulfate from the simple copper sulfides, andreacting said solution with H₂ S for precipitation of zinc valuestherefrom as zinc sulfide.
 2. A process as claimed in claim 1, reactingsaid solution of ferrous sulfate containing dissolved zinc values withH₂ S at a temperature of from ambient temperature to 100° C. with apartial pressure of H₂ S of from about 30 to about 100 psi.
 3. A processas claimed in claim 1, for separate recovery of any nickel and cobaltvalues contained in the thermally activated iron containing copper andzinc sulfides which comprises reacting the solution of ferrous sulfatecontaining dissolved zinc, nickel and cobalt values with H₂ S at atemperature of from ambient temperature to 100° C. with a partialpressure of H₂ S of from about 30 to about 100 psi for precipitation ofzinc sulfide, and raising the temperature of the solution up to 150° C.at a partial pressure of H₂ S of about 100 psi for precipitation ofcobalt and nickel sulfides.